The present invention relates to a process for leaching refractory sulphide and/or carbonaceous ores or concentrates and in particular is directed towards a method of recovering precious metals contained in the ores or concentrates.
Gold is generally extracted from gold containing ores by treatment with a cyanide solution which solubilizes the gold. However, in some ores, the gold is present as microparticles encapsulated within the ore. The gold in such ores cannot be extracted by traditional cyanidation techniques. These types of ores are known as refractory ores and are typically sulphide and/or carbonaceous ores. These ores may also contain, along with sulphides, other compounds of other Group VIA elements such as Selenium and Tellurium.
In order to extract gold from refractory sulfide ores, the ores must first be treated to liberate the gold so as to be accessible to cyanide leaching. A refractory ore is typically treated by oxidizing the ore which results in the chemical destruction of the refractory component of the ore, liberating precious metals for subsequent recovery. Known methods of oxidising refractory ores include roasting, bacterially assisted leaching and leaching the ore at elevated temperatures and pressure under acidic conditions.
An example of such a process for treating sulphide ores is known as the Sherritt process which includes the steps of feed preparation, pressure oxidation in the presence of acid and oxygen, solid/liquid separation, liquid neutralization and gold recovery from the oxidised solids by cyanidation. The operating conditions required in this process are at temperatures of between about 150xc2x0 C. to 210xc2x0 C., a total pressure of 2,100 kPa, a pulp density of 20% to 30% solids by mass, acid concentration of 20-100 g/L and a retention time of two to three hours. The oxidation must be carried out in an autoclave and requires a source of oxygen.
The capital costs associated with providing the autoclaves and meeting a high oxygen demand are high and may be prohibitive for construction at remote sites, for medium to small scale operations and for low grade ore. It is possible to leach under less aggressive conditions but in this case, the leaching rates and recovery are too low to be economically viable.
Attempts have been made to reduce the aggressive conditions and to lower the pressures and temperatures required whilst maintaining the economic viability of refractory ore treatment processes. For example, Australian patent application no. 27182/92 describes subjecting an ore concentrate to fine grinding prior to leaching. Fine grinding to an 80% passing size of 15 micron or less enabled the leaching to be carried out under less aggressive conditions at temperatures of 95-110xc2x0 C. and pressures of about 1000 kPa.
Thus, whilst some progress has been made in reducing the operating parameters when using oxygen as the oxidant, the leach must still be carried out under pressure.
U.S. Pat. No. 5,536,480 also describes subjecting an ore to fine grinding prior to acid pressure leaching. In this case, the refractory sulfide ore contains carbonaceous material and the ore is ground to a particle size of 40 microns or less. It was found that in order to obtain an acceptable gold recovery it was necessary to oxidize the material at a minimum temperature of 200xc2x0 C. and to obtain a minimum sulfur oxidation of 96%.
It is also known to oxidatively leach mineral species with ferric ions under acidic conditions. Ferric ion is typically a more effective oxidizing agent than oxygen which means that oxidation with ferric ions can be carried out under less aggressive conditions. Oxidative leaching at atmospheric pressure using ferric ions is known. A disadvantage of leaching with ferric ions is that the ferric ions are reduced to ferrous ions during the leaching reaction. As leaching solutions are recycled, therefore, ferric ions must be regenerated by oxidizing the reduced ferrous ions.
A further disadvantage with the aforementioned leaching processes for precious metal recovery is that they operate under acidic conditions. One difficulty with leaching under acidic conditions is that ores which contain an organic carbon fraction, known as carbonaceous ores cannot readily be processed to obtain acceptable levels of precious metal recovery. Precious metals such as gold are typically recovered from a leach solution by a cyanide leach stage. The reduction in recovery is due to absorption of the precious metals by the organics, during the cyanide stage. The absorbed metals cannot be recovered by cyanidation without a further pre-treatment stage designed to destroy the carbonaceous matter.
In order to address the problem of organic carbon, carbonaceous refractory ores are commonly treated by roasting to convert carbonaceous matter in the ores to gaseous carbon dioxide, or by treatment with a strong chemical oxidant, such as chlorine, to oxidize the carbonaceous matter. Both methods are expensive and are not economically viable for treatment of low grade materials.
Still further, as leaching steps such as cyanidation require alkaline conditions, the acid must be removed prior to cyanidation.
The problem of removing acid prior to cyanidation may be overcome by leaching under alkaline conditions. Leaching of nickel and cobalt under alkaline conditions using ammonia/ammonium salts is well known. However, a major disadvantage of alkaline leaching is that when iron containing ores such as pyrite are oxidised, the iron which is leached precipitates as a passive iron oxide layer on the mineral particle. This layer inhibits further oxidation with the result being that the extent of leaching under alkaline conditions is less than under acidic conditions. This translates to a lower recovery of precious metals.
Still further, alkaline leaching of refractory materials requires elevated pressure and temperatures and an oxidant for the leaching to occur. However, even under aggressive alkaline conditions, recovery of precious metals is often less than that for acid leaching. Further, base metals such as copper and zinc are insoluble at high pH. Thus, alkaline leaching is unsuitable for leaching ores or concentrates where recovery of base metals from base metal sulphides such as chalcocite, sphalerite or chalcopyrite is required. For these reasons, commercial and academic interest has been directed towards acid leaching.
Most of the literature relating to alkaline leaching is directed towards the use of water soluble alkalis such as sodium or potassium hydroxide and ammonia. A disadvantage with these reagents is that iron is precipitated primarily as jarosite. Jarosite inhibits gold recovery and is also an environmentally unacceptable residue. Also, hydroxide reagents and in particular sodium hydroxide are prohibitively expensive.
The use of cheaper alkalis such as lime has been proposed. However, to date, leaching of iron sulphide materials with lime has been unsuccessful in that leaching is incomplete and subsequent precious metal recovery is low. For example, an earlier study of alkaline oxidation of pyrite for gold recovery using lime achieved only 30 to 40% gold recovery which offered little improvement over direct cyanidation of the pyrite. This is believed to be due to passivation of the mineral by precipitation of a gypsum/iron oxide layer.
Limestone is another alkali which is relatively cheap. Limestone is typically used in the neutralization of acidic leachates. However, limestone is considered to be insufficiently reactive and/or soluble in alkaline systems to be able to be used for alkaline leaching.
From an economic point of view, it would be desirable to be able to leach refractory materials for precious metal recovery under mild alkaline conditions and using reagents other than the expensive hydroxides.
As mentioned above, it is known that the oxidation rate under acidic conditions can be increased by fine grinding to increase the surface area of the mineral particles. Such an increase may be predicted given that there is a larger surface area exposed to the oxidizing agents. However in the alkaline system, this effect is substantially reduced in view of the formation of the passive iron oxide layer on the particles. The rate determining factors in the alkaline systems are believed to relate to the formation of the passivating iron oxide layer and diffusion of oxygen through the layer. Thus, workers in the field have concentrated on increasing the extent of alkaline leaching by using strong, soluble alkalis, by modifying the leaching conditions so as to minimise formation of the passive layer and/or influence the diffusion rate through the layer.
One study suggests leaching at higher temperatures or at relatively concentrated solutions of reagents. The reason for this is to rapidly produce a passive layer which is unstable and subject to cracking. It is believed that at lower temperatures, the layers grow more slowly and are more stable. Another suggestion has been to use additives which may react to dissolve the layer or to make the layer more permeable.
However, to date no method has been proposed which is able to economically leach iron containing refractory ores and concentrates under alkaline conditions and which also enables good recovery of precious metals from the ore or concentrates.
The present invention is based on the surprising and unexpected discovery that leaching refractory sulfide and/or carbonaceous materials under alkaline conditions can be successfully achieved by careful selection of the particle size of the material to be leached. Even more surprising it has been discovered that not only can the leaching efficiency be improved but that leaching can also be successfully conducted under relatively mild conditions. Still further, it has also been discovered that the activity of any carbonaceous matter in the feed material can also be substantially reduced as part of the oxidation process, to a point where it will not compete with the activated carbon added to commercial gold/silver recovery processes, where cyanide is used to leach the precious metals.
According to a first broad form of the invention there is provided a method of processing a mineral composition comprising a refractory material the method comprising milling the composition to a particle size of P80 of less than 25 microns and leaching said composition with a solution comprising lime and/or limestone in the presence of an oxygen containing gas.
In the present specification and claims the term xe2x80x9crefractory materialxe2x80x9d includes refractory sulfides such as pyrite in which precious metals are encapsulated, ores which contain carbonaceous material and telluride or selenide materials. In the present specification and claims, the use of the term xe2x80x9corexe2x80x9d Includes not only ore per se but also includes concentrates, slimes, tailings, spoil and waste materials which may have a recoverable amount of precious metal values. Carbonaceous material refers to materials having an organic carbon fraction which may include graphite, bituminous or partly bituminous material.
The method of the present invention is particularly applicable to a composition including an iron containing refractory sulphide, selenide or telluride material or a mixture of such materials, with or without carbonaceous material present. Examples of such materials include pyrite, marcasite, arsenopyrite and arsenic bearing pyrite, troilite and pyrrhotite.
According to a further broad form of the invention there is provided a method of processing a mineral composition comprising an iron containing refractory material the method comprising milling the composition to a particle size of P80 of less than 25 microns and leaching said composition with a solution comprising lime and/or limestone in the presence of an oxygen containing gas.
Other non-iron or minor iron containing materials may also be present in the composition, examples of which include stibnite, tetrahedrite, argentopyrite, calaverite, altaite, gold bearing selenides, tennantite and pentlandite. The method of the present invention is also applicable to a composition including carbonaceous matter, where the carbonaceous matter would otherwise interfere with the precious metals recovery process. Suitably the composition would not include economic amounts of base metal sulphides containing copper or zinc. Generally a composition which includes appreciable amounts of these base metals would not be treated directly by the method of the present invention. The reason for this is purely economical as copper and zinc precipitate under alkaline conditions and thus cannot be recovered easily. Under acidic leaching conditions, copper and zinc are solubilized and can be recovered by conventional SX/EW techniques. Use of flotation or other separation technologies to produce a base metal concentrate and a separate refractory iron sulphide concentrate would be obvious to one skilled in the art as a way of treating these materials.
The method of the present invention is in particular directed towards the treatment of refractory materials containing precious metals such as gold, silver and platinum. Suitably, the refractory materials are in the form of flotation concentrates although the method is suitable for ores if the economics are favourable.
In the method of the present invention, the composition is finely ground to a particle size of 80% by mass passing less than 25 xcexcm. A typical particle size range is between 80% passing 2-25 xcexcm and preferably between about 80% passing 5-15 xcexcm.
A preferred apparatus for producing the finely ground material is a stirred ball mill. However, it will be appreciated that several other suitable types of comminution apparatus may also be used.
The leaching is carried out using lime and/or limestone as the alkali reagent. Lime, limestone or a mixture thereof may be used. Preferably a mixture in the range of 40-95% limestone is used. The lime and/or limestone is added in an amount such that the pH of the system is between about 6-12 and preferably about 6-9.
Typically about 100 to about 1200 kg, of lime and/or limestone is added per tonne of solids. The amount of lime and/or limestone which will need to added to maintain a desired pH will generally vary according to the amount (if any) of sulphuric acid produced by sulphide oxidation. Generally about 800 kg of lime and/or limestone would be added.
The present inventors have also surprisingly discovered that not only can the overall amount of leaching be increased but that such an increase can also be obtained by leaching under milder conditions than has hitherto been possible. Suitably the method of the present invention can be carried out at ambient pressure. This avoids the use of expensive pressure reactors and autoclave equipment. The preferred operating temperature is between about 50xc2x0 C. up to the boiling point of the mixture. Typically the maximum temperature is about 95xc2x0 C.
The leaching reaction can thus be carried out in open tank reactors. Excess heat is removed by evaporation of the solution. This avoids the need for costly heat exchangers. Heat can be easily introduced by known methods such as the injection of steam.
The leaching reaction is carried out in the presence of an oxygen containing gas. When the reaction is carried out in an open tank reactor the gas is typically introduced by sparging. The gas may be oxygen, air or oxygen enriched air. The gas flow is dependent upon the amount of oxygen required to sustain the leaching reaction and is typically about 0.01-0.5 vvm (vessel volumes per minute). The oxygen consumption of the process is typically between about 200-1000 kg of oxygen per tonne of solids.
After the composition has been leached, the mixture can be further treated by known methods to recover precious metals, principally by cyanide leaching.
Cyanidation occurs under alkaline conditions. As the leach solution is already alkaline, the solution may be subjected directly to cyanide leaching. If desired the slurry may be thickened prior to cyanide leaching. A further advantage of the current invention is that the type of compounds that form in the alkaline leaching system are not reactive toward cyanide, and will not consume high levels of cyanide in the gold recovery stage. Compounds formed under acidic leach conditions often consume significant amounts of cyanide in the gold recovery stage, increasing the process costs.
According to a further broad form of the invention there is provided a method of recovering precious metals from a mineral composition comprising a refractory material, the method comprising;
grinding the material to a particle size of 80% passing 25 xcexcm or less;
leaching the ground material in the presence of lime and/or limestone and an oxygen containing gas; and
subjecting the leached material to a further leaching step to recover any precious metals.